1. Field of the Invention
The present invention relates to recovering copper from a copper-bearing ore and more particularly to a hydrometallurgical process and system for recovering copper from chalcopyrite concentrate.
2. State of the Art
Because of increasingly stringent environmental controls, alternatives are being sought for the conventional methods for producing commercial grade copper. Many of the new methods under investigation are hydrometallurgical. That is, copper from a copper-bearing mineral is first dissolved into solution and then recovered from the solution. Typically, the copper is recovered by electrolytic techniques.
Chalcopyrite (CuFeS.sub.2) is the primary copper-bearing mineral mined in the world. In the past, the efficient hydrometallurgical recovery of copper from chalcopyrite has been hampered by two major obstacles. First, chalcopyrite concentrate is extremely difficult to leach. That is, it is extremely difficult to dissolve copper from the chalcopyrite concentrate into solution. Second, known chalcopyrite leaching techniques which are successful in efficiently leaching copper into solution also leach iron from the chalcopyrite into solution. Because excessive iron in solution interferes with the electrolytic recovery of copper, a separation of the dissolved copper and the dissolved iron must be accomplished before the copper may be recovered from the solution. For these reasons, and because of increasing environmental demands, a great deal of research has been conducted, both by academic and commercial interests, toward the efficient recovery of copper from chalcopyrite by hydrometallurgical techniques.
A number of investigations have been directed to hydrometallurgical techniques which involve the direct leaching of chalcopyrite in an electrolytic cell with the concomitant recovery of pure copper. Processing of the chalcopyrite in this manner results in the problem stated above, namely, iron dissolved from the chalcopyrite accumulates within the electrolyte contained in the recovery cell. Thus, complicated and expensive purification systems are required to remove iron from the electrolyte. The simultaneous leaching and electrowinning of copper from chalcopyrite in an electrolytic cell suffers from the additional basic problem that the cell generates only 50% of the oxidant required to leach copper from the chalcopyrite at the same rate at which it is being electrowon. The remaining 50% of the required oxidant must be supplied from an external source.
To avoid iron build-up in the electrolytic cell, methods have been developed for removing iron from the chalcopyrite processing circuit prior to introduction of the copper solids to the cell. According to one such method, a stoichiometric amount or more of sulphuric acid is utilized to dissolve both the iron and the copper contained in the chalcopyrite concentrate at high temperature and under high oxygen pressure. The copper-iron solution thus produced is then oxidized under autoclave conditions to precipitate the dissolved iron. The resulting copper solution is then processed in an electrolytic cell. According to a second well-known approach, the chalcopyrite concentrate is roasted prior to leaching to make the copper soluble and the iron insoluble.
Both of the above methods have serious associated problems which prevent them from becoming competitive with conventional methods for producing copper. The use of high temperature and high pressure autoclaves is expensive both in capital costs and in operation. In addition, autoclaves cause the oxidation of at least some of the sulfur contained in the chalcopyrite to sulfate, an undesirable end product which must be removed from the system. If roasting is used, the most of the pollution problems encountered by conventional methods are introduced and it becomes expensive and difficult to operate within government standards.
According to another approach to chalcopyrite processing, chalcopyrite leaching without the use of autoclave conditions is accomplished by initially fine grinding the chalcopyrite concentrate to a particle size of one micron or less. The finely ground chalcopyrite concentrate is then leached in a low acidity, ferric sulfate solution to produce a pregnant leach liquor containing cupric and ferrous ions and solid elemental sulfur. The pregnant leach liquor is then separated from the remaining solids including elemental sulfur and unreacted chalcopyrite concentrate. Copper is then recovered from the pregnant leach solution by one of two alternative methods. According to one method, the pregnant leach solution is treated for iron removal by precipitating ferrous sulfate. The copper-containing solution is then electrolized in a diaphragm cell to produce elemental copper. According to the second method, the pregnant leach solution is subjected to a conventional solvent extraction process to isolate the copper ions which are then reduced to elemental copper by electrolysis. The remaining solution, which contains ferrous iron as dissolved ferrous sulfate, is then treated with sulfuric acid and oxygen to oxidize the ferrous iron to ferric. The ferric iron is then recycled to the leach step for use in treating further amounts of ground chalopyrite concentrate. Such a process is taught by U.S. Pat. No. 4,115,221.
The process taught by U.S. Pat. No. 4,115,221 suffers from a number of disadvantages. To achieve an acceptable leach, the process requires that the chalocopyrite concentrate be ground to a maximum particle size diameter of one micron. Grinding to such a particle size requires a relatively large power input. For example, about four times the power is required to grind the concentrate to a one micron size than is required to grind it to the five micron size typically used in chalcopyrite processing. Another disadvantage of the process is that ferric iron for use in the leach is generated in a clarified solution with no chalcopyrite present. The maximum amount of ferric iron that can be retained in such a solution is about 60 grams per liter. This results in the leaching of only about 15 grams per liter of copper from the chalcopyrite into solution. This means that either dilute copper streams are produced or repeated cycling of the solution through oxidation and leach steps is required to increase the copper concentrations to a level suitable for efficient recovery. Furthermore, if ferrous sulfate is crystallized from a solution containing significant amount of copper, copper is crystallized with the iron. This results in a loss of copper with the iron or requires further processing steps to recover the copper.
According to another chalcopyrite leaching process, chalcopyrite is initially leached with a copper sulfate solution to form a slurry containing the insoluble copper sulfide, digenite, together with soluble iron sulfate and sulfuric acid. The digenite is separated from the iron sulfate solution and treated to produce elemental copper and sulfur. Alternatively, the products of the initial leach are subjected to a secondary leach in which the digenite is converted to a copper sulfate solution while iron sulfate is precipitated as jarosite. The copper sulfate solution is then separated from the remaining solids and elemental copper is recovered from the solution by electrolysis. The processing of the digenite by electrolysis requires only one-half of the oxidant per unit of copper leached as does processing the chalcopyrite concentrate. This means that an electrolytic cell supplies all of the required oxidant for the simultaneous leaching and electrowinning of the digenite. This process is taught by U.S. Pat. No. 3,957,602.
However, the process taught by U.S. Pat. No. 3,957,602 also suffers from a number of disadvantages. The conversion of chalcopyrite to digenite is carried on at temperatures greater than 100.degree. C. and preferably at temperatures of 180.degree.-200.degree. C. This requires the utilization of expensive autoclaving equipment. The process also generates significant amounts of excess free acid.